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By Jean-Marc Simard, Vincent Jourdain, Denis W. Roy

"Seven lithologic units, informally characterized, were recognized in the vicinity of the Montauban orebody. The first five lie in stratigraphic order. From structural bottom to structural top these are: (1) grey composite gneiss; (2) dark grey quartz-biotite gneiss; (3) quartz-biotite-garnet gneiss hosting several calcareous, aluminous and magnesian members and gold mineralization; (4) light grey quartz-muscovite-(sericite)-biotite gneiss; and (5) metabasafts. The last two intrude the sequence: (6) amphibolite dykes and (7) pegmatite veins. The orebody lies on the inverted limb of an overturned synform verging to the west with the metabasafts marking the core of the fold. As the foliation is folded, this fold is attributed to deformation ""D2""; deformation ""DI"" being responsible for the formation of the foliation. A late brittle deformation developed shear zones with probable strike-slip motions. The refined stratigraphy and the identification of the structural position of the orebody provide new tools for regional interpretation and prospecting.RESUMESept unités lithologiques ont été définies informeffement à proximité du gisement de Montauban. Les cinq premieres reposent en ordre stratigraphique; ce sont de la base au sommet structural: (1) le ""gneiss mixte gris""; (2) le gneiss gris forcé à quartz biotite; (3) le gneiss a quartz-biotite-grenat comprenant plusieurs membres calcareux, alumineux et magnésiens ainsi que la minéralisation aurifère; (4) le gneiss gris clair à quartz muscovite- Isericitel-biotite; et (5) les métabasaltes. Les deux autres unites sont intrusives dans la sequence: (6) les dykes d ‘amphibolite et (7) les veines de pegmatite. Le gisement se sitlie sur le flanc inverse d'un synforme déverse vers I' ouest dont les métabasaltes occuperaient le cœur. Comm e la foliation est plissée, ce pli est considère appartenir d la deformation ""D2""; la deformat ion ""DI"" etant responsable de la formation de lafoliation. Une deformation cassante tardive se traduit par des zones de cisaillement, d rejet apparemment directionnel. Cette stratigraphie plus précisé et l'indentification de la position structurale du gisement fournissent de nouveaux omits de prospection et d'interprétation regionalé."

Jan 1, 1987

By K. Schrijver, G. Beaudoin

"Four occurrences of galena-cemented sandstones, exposed in the Lower St. Lawrence - Gaspésie area, are described. They constitute a new, insufficiently explored type of deposit in Quebec.One mineralized body, a minor occurrence, is hosted in Lower Devonian rocks deposited in an estuarine, intertidal river channel during a marine regression; major cements are dolomite and sulphides. Another occurrence, a drilled prospect, is hosted in Upper Silurian rocks deposited unconformably on Cambro Ordovician basement during a marine transgression. With galena and quartz as main cementing minerals, it is similar to some well-known, economic sandstone-lead deposits. A third and fourth, a drilled prospect and a minor occurrence, are hosted in Upper Cambrian rocks deposited on continental slopes. In both, sulphides mainly occupy a secondary porosity. The differences between the third and fourth occurrences are: (1) cementing translucent minerals (carbonates and barite in one; chlorite in the other); (2) chronology of final subsurface dissolution (prior and subsequent to sulphide precipitation, respectively); and (3) important fracture porosity only in the barite-bearing deposit.The diversity of the four occurrences, and the unequal geological knowledge of their respective features, prohibit the formulation of uniform guidelines for further exploration.IntroductionThe attention of the mineral exploration community is drawn to four occurrences of galena-cemented sandstones (Fig. 2A) in the Cambrian of the Taconian Orogen and in the Siluro-Devonian basin of the Lower St. Lawrence and Gaspésie area (Fig. 1). In view of the fact that sandstone-lead deposits (Bjerlykke and Sangster, 1981) are, or have been, mined in several countries (e.g . Laisvall in Sweden: see Rickard et al., 1979; Largentiere in France: see Foglierini et al., 1980), the mere presence of occurrences in Quebec indicates that the mineral potential of some parts of the Appalachians requires further evaluation.Exploration for the deposits considered here is essentially restricted to geological methods and drilJing (e.g . Mainguet Suares and Samama, 1969), since the response to geophysical methods is virtually nil, at least in the initial, surficial stage of an exploration program. As the particular disciplines of geology utilized in exploration depend a great deal on the geological setting of a mineral occurrence, the diversity of the Quebec deposits is of prime importance. Therefore, the four occurrences (Fig. I, Table 1) studied by the authors are herein briefly described with emphasis on their stratigraphic and tectonic setting, and on the evolution of pore space and sequential occlusion of pores and fractures. Prior to further research in and around the occurrences, now in progress (e.g. Beaudoin, 1987; Beaudoin et al., 1987), it is hoped that the following outline will be useful to the explorationist."

Jan 1, 1987

By R. G. Burn

"A brief outline will be given of the concept of risk and its implications with regard to investment in exploration. The relationship between risk and decisions under uncertainty will be examined as a prelude to identifying those factors which contribute to exploration risks:consideration of the effects of corporate objectives (e.g. survival, expansion, diversification) on exploration strategy and ultimately on success or failure, exploration and mining as an investment (the relationship between risk and reward) and the desired return on investment; exploration odds, funding levels, staying power, financial exposure and the concept of gambler's ruin;origins of uncertainty in exploration decision-making-man-made risks and natural hazards; the nature of introduced constraints at each exploration decision stage and their effect on the rate of success;improving the chances of success and the invest reduction of risk through sound exploration strategy, effective project generation and efficient target detection capabilities."

Jan 1, 1984

By Don Thompson

It isn?t a lie if you don?t tell the whole truth The pictures National Geographic didn?t print Other photos National Geographic didn?t print?

May 1, 2009

By H. A. Bilgin, G. G. U. Aldas

"Ground vibration is the natural result of the blasting process. Many empirical formulae and methodologies including analysis of travel path effects have been developed. The scaled distance, which is the most popular approach for the calculation of the attenuation of particle velocity in the ground, has led to significant advances in the prediction of particle velocity when both explosive charge and distance vary. However, this approach does not consider the effect of ground anisotropy and existence of structural geological features on ground vibration. Therefore, in this study, the effects of these rock mass properties on types, amplitudes, and frequencies of ground vibration waves at four surface mines are examined. The results, which were implemented at Kangal coal mine, Bastas quarry and Kíbrís quarry, provided guidance for the extensive study at Orhaneli surface coal mine.As a result of the study, the authors came to the conclusion that vibration amplitudes change depending on the ground on which sensors were deployed and the conditions depending on whether measurements were done near major structural geological features. It was determined that the vibration amplitudes, measured at equal distances from the source but in different directions, change because of the propagation velocity anisotropy. Besides the change of wave type due to the kind of ground at the measurement location, it was determined that vibration anomolies were due to structural features."

Jan 1, 2004

By M. Sharafodin, A. A. Hassanipak

Abstract - The essential aims of additional borehole drilling are to improve the reliability of grade and tonnage estimates in each reserve class and to increase ore tonnages. The "GET" function presented in this paper considers strategies for achieving both of these goals simultaneously, and therefore is advantageous for selecting sites for additional boreholes. The "GET" function is either a linear or a non-linear product of three variables G, E, and T: f(G,E,T,) = G??E??T??where the values of any or all of the exponents ?, ?, and ??may differ from unity at the discretion of the user. G and E are the average estimated block grade and the average estimation error for ore blocks in one vertical column, and T is the compounded ore thickness within the column. To illustrate its utility, the GET function has been used for determination of the most advantageous sites for additional drilling in the Shah-Kuh Pb-Zn deposit in west central Iran.

Jan 1, 2004

By Tarrant Elkington

Underground mine planning practices are typically manual and time consuming and will often produce sub-optimal solutions due to the application of generic rule-of-thumb assumptions or ?tried and tested? choices. The development of computer-automated tools for underground mine planning enable a wide range of scenarios to be considered rapidly in arriving at an optimised solution, subject to input assumptions. We have applied these tools to demonstrate ?optimised? solutions that challenge industry standard practices. Further, we have incorporated grade and price/cost uncertainty into these tools to gain an understanding on how the consideration of uncertainty may further alter strategy selection. Finally, we enable the tools to dynamically adjust cut-off grades in conjunction with the evolution of prices/cost.

May 1, 2010

By Panagiota Tsolis-Katagas, Basilios Tsikouras, Tassos A. Grammatikopoulos, C. Katagas

"Mineralogical techniques including microscopic methods, reflected and transmitted light microscopy, Cathodoluminescence (CL) and Scanning Electron Microscopy equipped with an Energy Dispersive Spectrometer (SEM-EDS), are useful in determining the quality of industrial minerals. They are essential for studying ores and metallurgical products including the evaluation of physical characteristics of the mineral, paragenetic relationships, associations with possible deleterious contaminants (e.g., sulphides), spatial distribution within the ore zones and liberation characteristics. The methods are inexpensive, and provide quick identification and distinction among the industrial minerals and the gangue minerals in a section. Additionally, SEM is necessary to reveal complex textural sequences and 3-D measurements of minerals and quick semi-quantitative analysis.Common analytical methods include Electron Microprobe Analyses (EMPA) and Whole Rock Analysis (WRA) (commonly with XRF and ICP) for major and trace elements. The EMPA analyses of industrial minerals provide information on the purity of the analyzed grains as opposed to the bulk chemical analyses that may be misleading due to contamination by other associated common “gangue” minerals. These techniques have been implemented for the case studies of wollastonite, V-magnetite and muscovite. Therefore, these methods can provide extremely useful information and lead to the implementation of a successful pilot-testing programme.MINERALOGICAL EVALUATION OF DEPOSITSDespite their considerable and increasing economic importance, industrial mineral deposits have traditionally been neglected by economic geologists, and are only rarely the subject of the detailed analysis widely lavished on metallic mineralization. Moreover, in that they generally represent concentrations of “normal” rock-forming minerals, albeit commonly with particular textural and/or geochemical features, they have tended to fall between the mainstream of petrogenetic studies and mineral deposit research (Grammatikopoulos 1999). For many industrial minerals, little attention has been paid to the factors responsible for their local concentration as large and high-grade bodies. This comparative neglect has particularly been the case for minerals such as talc and wollastonite, widespread in metamorphic rocks, in which they are reasonably interpreted as components of assemblages derived, respectively, from pelitic sediments, ultramafic igneous rocks and siliceous limestones."

Jan 1, 2003

By D. Imeson, T. Grammatikopoulos, M. Aghamirian, M. Oliazadeh

"More than 70% of potash is produced worldwide by froth flotation. Flotation of potash was developed in the early 1930s. Amines are widely used for flotation of sylvite (KCl) from halite (NaCl) and insoluble materials such as clays. Amine collector is also absorbed by clays. Therefore, clays are usually separated by scrubbing and de-sliming prior to flotation. Alternatively, slimes can be floated ahead of potash flotation. Flotation recoveries of sylvite are commonly in the range of 85%, while losses can occur in slimes, flotation tailings and make-up of process brine. Ores containing free coarse-grained sylvite are generally screened, whereas both fine and coarse fractions are conditioned and floated separately. The unique features of potash flotation include solubility of sylvite in water, the size of particles floated and fast kinetics. Particles up to 2.3 mm can be floated and about 50% of flotation feed can be removed in the rougher stage in few minutes. The main challenges in potash flotation are the recovery of potash from slimes, flotation efficiency in the presence of carnallite and recovery of coarse potash grains. A review of potash flotation shows that there is still room for improvement in both coarse and fine potash flotation. The present paper reviews the most significant flotation parameters including reagent type and dosage, flotation circuits, flotation equipment size and new developments in potash flotation. The latter include new reagents, flotation technologies (Hydrofloat, Jameson and column cells) larger flotation cells and slimes flotation.INTRODUCTION Potash is the common industrial term for potassium in various chemical combinations and minerals soluble in water. About 90% of potash produced worldwide is used as agricultural fertilizers. Flotation is the main treatment process of sylvinite (KCl-NaCl) ores for recovery of sylvite (KCl). Potash concentrate should contain a minimum of 95% KCl and be relatively free from 0.2 mm KCl fines. The flotation process for beneficiation of potash ores was developed in United States and Russia concurrently in the early 1930s. In 1934, Potash Corporation of America applied flotation process in a commercial scale (Tippin R.B. and Huiatt J.L). In the initial application, halite (NaCl) was selectively floated by employing lead chloride as an activator and soap as collector. The sylvite concentrate was reported in the cell underflow."

Jan 1, 2012

By E. Roman, M. Brissette

"For its end of life, Kemess Mine processed its difficult-to-treat hypogene ore from the east pit, which is a lower grade copper and higher pyrite content gold ore. Metallurgical testwork showed that improvement was possible with finer grinding. In July 2010, small grinding media were introduced into the regrind mill resulting to a nearly 50% decrease of the power draw with a finer grind, an increase of 1.1% in the final copper grade, an increase of 1.2% in copper and 2.9% in gold recoveries at the cleaners. The overall recoveries of copper and gold increased by 0.6% and 1.3% respectively.INTRODUCTION Kemess Mine was located at 430 km North West of Prince George in North Central British Colombia. Its production started in May 1998 by Royal Oak Mines and was acquired by Northgate Exploration in 2000. Kemess reached its end of life production with its last of production on March 9th, 2011. The concentrator nominal capacity was 56,000 TPD consisting of a primary gyratory crusher feeding two identical conventional SAB circuits totalizing 35,840 kW available for grinding. Rougher flotation consists of four parallel rows of eight 130 m³ tank cells to produce a concentrate of 2% Cu and 6 gr/t Au which was pumped to a 5 m x 10.8 m regrind ball mill. The regrind installed power was 4,480 kW to generate a final grind of 80% passing 40 microns. This product fed the two stage cleaner flotation circuit, first stage being two parallel rows conventional 17 m³ Denver cells and second stage, two 3.3 m x 11.0 m flotation columns in series, designed and constructed by Kemess, to produce a 22% copper concentrate grade. The cleaner flotation middlings were returned back to the regrind mill (Froehling, Mohns, Roman & Grady, 2005; Froehling, Mohns, Rajwani & Fullam, 2007). The large copper-gold porphyry deposit was composed of 15% of oxidized supergene and leach cap ore and 85% of hypogene ore having three major alteration zones, mainly characterized by their sulphur/sulphide content. Chalcopyrite is the main copper mineral species with minor bornite. Gold was mainly present as free gold or electrum. The first alteration zone consisted of a phyllic alteration characterized by moderate and intense sericite alteration which had the highest mill throughput and twice the mine grade when in contact with the supergene horizon. The second zone consisted of a potassic alteration characterized by more competent groundmass and moderate to intense quartz stockwork which had a slightly lower mill throughput. Finally, a third alteration found beneath and on the east side of the deposit having almost no quartz stockwork and hardened groundmass, sulphide content is two to four times more. This zone had the lower mill throughput with significant drop in the mine average gold and copper grade. Through its mine life, the average life of mine copper head grade was 0.21% Cu and 0.64 gr/t Au. Head grades began to decrease from 2008 to be substantially lower in 2011 as shown in Table 1."

Jan 1, 2012

By K. Lim, J. Sobil, L. Galarza Castillo, F. Tozlu

"For mining operations globally, the scale and criticality of the grinding mill drive has meant that selection of the correct mill drive technology is paramount. With so much depending on the successful implementation of the electrical drive system— safety operation, timely installation, availability, energy efficiency, minimum maintenance and extend its use as tool to assist during mill liner exchange—customers are justified in their scrutiny and slow adoption of new drive technology. From the first installation of low, fixed speed synchronous motors to the introduction of the current source and further voltage source variable frequency drive, innovations require long periods of time and mining customers willing to push the industry forward. In recent years, the variable frequency drive has been the stepping stone for a new generation of mill drive technology. In use for decades in other heavy industrial applications, the drives allow for system optimization in a myriad of ways. The most dramatic of changes is in the design of the motor. Where in the past motor design was bound by line frequency, today the designer can optimize machine size and, more importantly, eliminate the need entirely for the synchronous motor. Induction motor design has now progressed to the point where low speed operation is possible at efficiency levels nearing those of the synchronous design. Moreover, the simplicity of the squirrel cage design means that availability, reliability and spare part requirements are greatly improved. As the next step in the evolution of large pinion driven grinding mills, the low speed induction drive offers a new reliable and robust system to the market.IntroductionAs deposit grades continue to decline, new ways of processing ever greater concentrator throughputs have been required to maintain throughput efficient and profitable operations. Furthermore, the high risk nature of mining developments requires a constant search for safer, higher availability and more efficient equipment design. Limitation of downtime is a key factor in ensuring mining operations can be sustained through volatile market conditions. Often processing the entire mine throughput through just 1 or 2 lines, the crushing and grinding circuits are of particular criticality to the uptime and, thus, profitability of any mining operation."

Jan 1, 2016

By Paolo Toscano, Dave Hall

"This paper discusses the project history, process design and layout of the Rainy River Gold Project mill. The circuit consists of a 55 x 83 primary crusher, 15 MW SAG mill and 15 MW ball mill. Both mills are twin pinion with LSS and VFD drives. A pebble crusher is located in the process plant with bypass and stockpile options. Gravity concentration followed by 30 hours of leach residence, carousel CIP, carbon recovery, cyanide destruction and tails disposal in a 1,428 hectare tailings management area (TMA).The Rainy River Gold Project is located approximately 65 km northwest of Fort Frances in north western Ontario. The project is scheduled to be commissioned and reach commercial production in mid-2017. Expected cost is $US877M and is comprised of an open pit mine, process plant, tailings management area and supporting infrastructure. The process plant is designed to produce gold/silver Dore’ bars.The project has been developed on the following timeline:• 1967 – First exploration• 1994 – 17 zone discovered• 2004 – 433 Zone discovered• 2005 – Rainy River Resources Ltd. purchases property • 2006 - ODM zone discovered• 2007 - Cap Zone discovered• 2008 - First NI-43-101 Resource completed• 2012 - PEA completed & Intrepid Zone discovered• 2013 - New Gold acquires Rainy River Resources• 2014 - Updated feasibility study completed & Environmental Assessment report submitted• 2015 - Detailed Engineering >95% complete, EA approved and construction commencesThe property is accessed by provincial highway 600 and is within 17 km of 230 kV power line and railway sidings. Key project criteria are summarized as follows:"

Jan 1, 2016

By Kazutoshi Haga, Batnasan Altansukh, Atsushi Shibayama

In this study, the dissolution behavior of copper and arsenic from five different copper concentrates containing 21-34.7 %wt. of Cu, 1.3-7.3 %wt. of As and 15.6-27.5 %wt. of Fe were investigated under high pressure oxidative leaching conditions. Leaching experiments were conducted using different leaching reagents such as ferric sulfate, sodium chloride, deionized water, sulfuric acid and ferric sulfate-sodium chloride mixture in order to compare the effectiveness of each on the dissolution of copper and arsenic from copper concentrates. The effects of ferric sulfate (0.1-1 mol/L), sodium chloride (0.1-0.5 mol/L) and leaching temperature (100-180C) were investigated under 1 MPa autoclave pressure , particle size >45 m, pulp density of 100 g/L, stirring speed of 750 rpm and leaching time of 60 min. Results showed that the most effective leaching reagents for the dissolution of copper are 0.34 M NaCl and a mixture of 0.1 M Fe2(SO4)3 - 0.34 M NaCl, in which As extraction did not exceed 7%. The copper concentration in the pregnant leach solution obtained after leaching using 0.1 M Fe2(SO4)3 - 0.34 M NaCl was 13-30 g/L, while the concentration of As and Fe were in a range between 0.3-1 g/L and 0.2-10 g/L at 160C. INTRODUCTION Copper sulfide minerals, such as chalcopyrite (CuFeS2), chalcocite (Cu2S) and covellite (CuS) are used for the production of copper in mining and metallurgical industries worldwide. The conventional pyrometallurgical processing followed by electrorefining is a common process for the production of high purity copper (Moskalyk, & Alfantazi, 2003; Padilla, Aracena, & Ruiz, 2012). However, the degradation of copper content and an increase in the amount of metal impurities, especially arsenic (As), antimony (Sb) and bismuth (Bi) in copper concentrates cause serious problems to the production of copper and environmental pollution (Basha, Selvi, Ramasamy, & Chellammal, 2008; Long, Peng, & Bradshaw, 2012). Hydrometallurgical processes are used to recover metals from refractory ores, concentrates and tailings by leaching (Long, Peng, & Bradshaw, 2012; Masloboev, Seleznev, Svetlov, & Makarov, 2018). During the last decades, hydrometallurgical processing for the recovery of copper from enargite-bearing copper ores and concentrates has been extensively investigated due to its relatively more environmentally benign and less capital costs than the conventional pyrometallurgical processes (Long, Peng, & Bradshaw, 2012; Masloboev, Seleznev, Svetlov, & Makarov, 2018; Choubey, Lee, Kim, & Kim 2018). However, several issues facing the hydrometallurgical processing are incomplete extraction of copper associated with the coprecipitation of iron compounds, difficult stabilization of toxic impurities and production of a contaminated copper product (Dreisinger, 2003; Safarzadeh, Moats, & Miller, 2014). Hence, a number of studies have been conducted on the selective dissolution of copper from copper ores and concentrates or selective removal of arsenic leaving a clean copper residue using various leaching reagents or lixiviants (Safarzadeh, Moats, & Miller, 2014; Mihajlovic, Strbac, Zivkovic, Kovacevic, & Stehernic, 2007; Tongamp, Takasaki, & Shibayama, 2009). More extensive studies on the selective extraction of copper from enargite- bearing copper ores and concentrates are much needed to improve copper recovery.

Jan 1, 2019

By S. Bastola, Y. P. Chugh

"This research determines shear strength and stiffness properties of bedding planes and joints in the roof strata rocks within 7.5 to 9 m (25-30 ft.) overlying the Illinois #6 coal seam (within the pressure arch zone). A shear test loading device was designed into a 1.334 kN (150 t) Forney Compression Loading Machine to perform direct shear tests in accordance with ASTM D5607-08 “Standard Test Method for Performing Laboratory Direct Shear Strength Tests of Rock Specimens under Constant Normal Force”. The data were used to determine peak and residual friction angles and the dilation angle. A total of 49 bedding plane samples were tested, out of which 46 (36 intact and 10 relatively weak and loose) samples passed QA/QC procedures in accordance to ASTM D5607. Samples from eight (8) different bedding planes- shale/limestone (SL), shale/sandstone (SSs), shale/bone (SB), laminated sandstone (LS), shale/shale (SS), bone/bone (BB), bone/limestone (BL), and limestone/limestone (LL) were tested. The number of samples tested for each bedding plane were: SL- 11, SSs- 8, SB- 5, BB- 4, LS- 6, SS- 9, BL- 1, and LL- 2. Peak and residual shear strength values at 2.75 MPa normal stress range 1.06 – 6.26 MPa and 0.82 MPa to 0.82 – 4.14 MPa, respectively. The average normal and shear stiffness values are about 11.98 GPa/m and 3.11 GPa/m. Dilation angles are low (<10?) and negative in some cases. Angle of sliding friction values range from 9? to 42?. Joint Roughness Coefficient (JRC) values were less than 10.INTRODUCTIONRock mass may be defined as the total in situ rock medium containing bedding planes, faults, joints, folds and other structural features. Therefore, it is discontinuous and may have heterogeneous and anisotropic engineering properties. Upon creation of an excavation in a rock mass, bedding planes and joints are the most likely points of failure initiation. The failure typically occurs either in tensile or shear mode. Upon failure initiation stresses are transferred to the neighboring rock mass and the failure progresses until stability is achieved through equilibrium of stresses and strength. An example of this mechanism is development of the pressure arch. This study determined the shear strength and stiffness properties in shear and compression of bedding planes and joints in the immediate roof strata rocks within 7.5-9 m ( 25-30 ft.) overlying the Illinois No. 6 coal seam (within the pressure arch zone). Since such data are currently not available for Illinois Coal Basin mines, researchers and engineers have been using estimated values for design purposes. The immediate roof lithology in the proposed mining area consists of shale 0.62 m; Bone 0.09 m (0.3 ft.); Limestone, 0.10 m (0.33 ft.); Shale 2.37 m (7.9 ft.); and sandstone 0.55 m (1.83 ft.). The bedding plane samples were provided by a mine in south-eastern Illinois operating at a depth of about 180 m (600 ft.)."

Jan 1, 2015

By Anil K. Beersing

Effective drainage systems are required for reclamation of mine-disturbed areas to control surface water runoff and avoid excessive erosion. For the conventional structural design approach, armouring material necessary to satisfy the design criteria should not move during the design flood event since displacement would uncover the erodible underlying materials. This paper discusses an alternative geomorphic approach for designing drainage channels and watershed landscape. The approach is based on quantitative studies of geomorphic processes operating in a watershed. A design based on the geomorphic approach replicates the configuration, channel regime, sediment volume and hydrology of natural and mature drainage systems. Movement of some armouring material is acceptable in such an approach because the channels are allowed to mature by natural degradation and aggradation, with erosion rates comparable to the natural environment. The key advantages of a geomorphic approach over a purely structural approach include reduced sizes of armouring material, geomorphologically sustainable landscapes, ?walk-away? or maintenance-free design and reduced liability in the future. The geomorphic approach is illustrated using the example of a coal mining operation in northwest U.S.A. The paper discusses field data collection, design criteria, typical design details, and construction methods.

Apr 1, 2005

By A. R. Miller, G. Desrivières, M. C. Bétournay, D. L. Barnes, P. Laliberté

"Over the last four years, an international initiative, championed by Vehicle Projects LLC, the Fuelcell Propulsion Institute and Natural Resources Canada, has been active in carrying out research projects to prove out the concept of applying fuel cell technology to underground mining vehicles. These have been carried out by a consortium consisting of mining companies, equipment manufacturers, technology developers, national laboratories and consultants, with the participation of Canadian and U.S. mine regulatory agencies.This clean technology, which produces only water and electricity, would allow several advantages to be obtained, in the health area (by eliminating diesel emissions and reducing noise generation), in lowering production costs (such as reducing ventilation required) and lowering greenhouse gas emissions.This paper outlines the results obtained in several key areas, such as, the operation of fuel cells in underground conditions, risk issues related to underground and surface operation of hydrogen vehicles and hydrogen production and delivery systems, and the design and mining performance of the world’s first hydrogen mining vehicle, a mine production locomotive. The work to date on the fuel cell loader, designing the power plant and power transmission systems, operational requirements, and operating cost reductions is also presented."

Jan 1, 2003

By W R. Honeywell

Two series of laboratory flotation tests were carried out to study the treatment of Faraday ore by two-stage flotation. First the acid-consumers and then the uranium minerals were floated. On ore ground in the Faraday mill grinding circuit, and using Faraday mill water in the flotation step, a recovery in the flotation concentrate of from 91 to 93 percent of the uranium was obtained in 45 to 50 per cent of the weight at a grade of 0.24 per cent U30s. The saving of acid in leaching the uranium-bearing flotation concentrate, as compared to leaching the whole ore, ranged from 30 to 46 per cent. The average over-all extraction resulting from flotation and leaching of the flotation concentrate was 88.1 per cent. The best collector for the acid-consumers was sodium oleate. Acintol F A-1 emulsion was the best collector for the uranium minerals, and, in the uranium flotation step, an acid pH in the slurry was found to be necessary. The reagent costs were estimated to be 22 to 26 cents per ton of ore.

Jan 1, 1963

By M. Vaillé, B. Bossé, N. Durupt, A. Michaut

Imouraren is a sedimentary uranium deposit (> 150,000 tU @ 800 ppm), located approximately 100 km North-Northwest of Agadez (Niger); its depth reaches 110 to 170 meters. For low-grades ores, heap leaching is commonly considered as the appropriate technology for metal recovery. More precisely, the ore is crushed (≤ 300 mm) and agglomerated before being stacked. In this case, sulfuric acid can be added during agglomeration and/or irrigation (for three months to a year) to improve uranium recovery. It is known that mineralogical, chemical and physical properties of the uranium ore impact recovery, yet many challenges remain to increase recovery using this process. For example, ore heterogeneity and spatial variability (e.g., variation in uranium, carbonate and/or clay contents) can affect heap leaching performances (i.e., metal recovery, acid consumption or neutralization, percolation rate). However, laboratory column tests using composite ore sample that do not consider ore heterogeneity are widely used to anticipate recovery. Consequently, a new approach that consists in using Imouraren drill core ore samples was recently investigated. This latter also enabled a better assessment of the main parameters that influence uranium recovery and proposed a new predictive model to optimize mining operations.

Jan 1, 2020

By P. R. A. Andrews, R. K. Collings

"There are more than 300 occurrences of fluorite in Canada. Deposits of economic significance, however, are confined to Newfoundland and Nova Scotia in eastern Canada, Ontario in central Canada, and British Columbia in western Canada. Fluorite has been mined and processed intermittently in Ontario and Newfoundland until November 1990, when Canada&apos;s only producer Minworth Ltd., a U.K. based company, with a mine and concentrator near St. Lawrence on the Burin Peninsula, Newfoundland, closed its operation.CANMET has processed ore samples from each of the major mineralized areas, and thirty-one studies were conducted between 1920and 1970. A summary of the important aspects of the various testwork studies is presented. The production of acid-grade fluorite was possible by flotation, and ceramic-grade, by flotation and gravity concentration methods. Gravity concentration methods included heavy media separation, jigging and tabling. Flotation was by direct and reverse techniques; associated minerals, quartz, calcite, barite, etc. were depressed by a variety of depressants. The function of the major depressants is discussed briefly. The various uses offluorite which include hydrofluoric acid, chlorofluorocarbons, aluminum production, petroleum alkylation, fluxing and ceramic applications etc. are discussed briefly."

Jan 1, 1993

By Bruce C. Smart

"Very little attention was given to the regulation of uranium mills by the Atomic Energy Control Board prior to 1974. The ore grade treated by Canadian mills in the past was in the order of 0.08 - 0.3% U and it was not difficult to keep the exposure of workers well below regulatory limits. Furthermore, uranium milling has also been substantially free of occupational illness because environmental control has been generally satisfactory (l).The discovery of higher grade ore bodies in Canada has resulted in higher radiation levels in one mill and this has partly contributed to a need for increased radiation monitoring in the mill. In addition, as greater insight is gained on the effects of radiation, the implementation of the ALARA principle is being stressed at all mills.The concept of ALARA was emphasized by the ICRP (International Committee for Radiological Protection) in 1977 and stands for '""as low as reasonably achievable'"". It is based on the premise that no proven threshold for potential effects of radiation exists and that the actual individual exposure should therefore be kept as far below the regulatory limit as possible, economic and social factors being taken into account."

Jan 1, 1982